This invention relates to the treatment of metal sulphide concentrates.
Oxidative roasting of pyrite (FeS2) is a standard way of producing sulphuric acid. Roasting is used on an industrial scale, e.g. for the production of zinc, copper, and nickel, even tin, molybdenum, and antimony and, in many cases, takes place in conjunction with one or more leaching or smelting operations. Sulphide roasting is used to oxidize some (or all) of the sulphur. The resulting SO2 is treated further, most commonly producing sulphuric acid. Other options for recovery of sulphur include the production of elemental sulphur, or liquid SO2.
Modem roasting processes usually use fluidized-bed reactors, which are energy-efficient, and have a high productivity because of their favourable kinetic reaction conditions. The SO2 content in the off-gas is typically 8 to 15% by volume.
For pyrometallurgical processing, the usual purpose of roasting is to decrease the sulphur content to an optimum level for smelting to a matte. Partial (oxidizing) roasting is accomplished by controlling the access of air to the concentrate; a predetermined amount of sulphur is removed and, for example in the recovery of copper, only part of the iron sulphide is oxidized, leaving the copper sulphide (for example) relatively unchanged. Total, or dead, roasting involves the complete oxidation of all sulphides, usually for a subsequent reduction process.
There are many modem pyrometallurgical processes in which roasting is not a separate step, but is combined with matte smelting. Flash furnaces employ sulphide concentrate burners that both oxidize and melt the feed, and are used extensively in the copper industry. Autogenous bath smelting is another alternative. Here a lance blows air or oxygen, together with concentrates and reductant, into a molten bath, and the energy released by the oxidation of the sulphur provides much of the required energy for the smelting process.
The roasting process has several effects:
a) Drying the concentrates
b) Oxidizing a part of the iron present
c) Decreasing the sulphur content by oxidation
d) Partially removing volatile impurities, for example arsenic
e) Preheating the calcined feed with added fluxes (for example, silica or limestone), in order to lower the energy requirement of the downstream process
Environmental concerns have highlighted the need to lower the emissions of sulphur from smelters treating sulphidic raw materials. These emissions emanate primarily from the furnaces and converters, either as fugitive emissions or as process gases vented up a stack. It should be noted that the typical 1 to 2% SO2 in the off-gas from reverberatory furnaces (for example) is too low for effective acid production.
The general trend in recent years has been to eliminate as much as possible of the iron sulphides (usually pyrrhotite) during the milling and flotation stages, in order to minimize the sulphur input to smelters.
Dead roasting, i.e. close to 100% sulphur removal, has the benefit of removing essentially all the sulphur at the beginning of a smelting process. Furthermore, in comparison with the intermittent nature of SO2 produced in a converting operation, a steady and almost optimum SO2 content of off-gas from a roaster requires a smaller and less expensive acid plant.
Various roasting techniques in the recovery of copper are described in the literature. 1-16 
In the Brixlegg process, copper was produced by electric smelting of dead-roasted chalcopyrite concentrate in a circular AC (alternating current) submerged-arc furnace, using coal as a reductant.
Brixlegg reports a 95% recovery of copper to blister, and levels of copper in the slag of less than 1% have been claimed. The crude copper averaged only 95% copper, and the operation has been discontinued1. Disadvantages of this process are the relatively high copper losses in slag, and the high electrical energy consumption.
An undesirable aspect of the Brixlegg process is the fact that lead passes into the final copper anodes and makes them fragile if the concentration is too high. On the other hand, the exceptionally high recovery of other metals related to copper makes the process of particular interest for treating ores which contain nickel and noble metals. (The nickel can be separated from the anode mud.)
A submerged-arc furnace has been used for treating dead-roasted calcine in a process developed by the US Bureau of Mines14, as was also used in the Brixlegg process. It was found that in order to produce a high-purity blister (2.2% total impurities) and low-copper-content slag in a submerged-arc furnace, a two-cycle procedure was required. Using this rather inconvenient and non-continuous procedure, recoveries as high as 98% were attained.
In the nickel industry, Falconbridge17-24 and Inco25-29 have worked on processes involving the smelting of roasted sulphide concentrates. These processes use six-in-line furnaces, commonly employed in that industry, which generally operate at temperatures around 1400xc2x0 C. The reduction reactions needed to provide the appropriate conditions for recovering metals from the oxides tend to raise the operating temperature of these furnaces. Consequently, large volumes of air are drawn into the furnace to cool the freeboard space of the furnace. This tends to result in high losses of the feed materials as dust. Dust losses of up to 25% of the feed have been mentioned20.
Nickel production has however been accompanied by a level of SO2 generation which is environmentally unacceptable. It has been recognised that a major means to reduce SO2 emissions is to increase the degree of sulphur elimination in the fluidized-bed roasters. However, the existing furnace technology is limited in the degree to which highly roasted concentrates can be handled. The higher degree of roast demands more strongly reducing conditions in the furnace to smelt more oxidized calcine feed, and to counteract slag losses. Higher coke addition rates are needed. Extra energy is generated by the additional coke combustion products, resulting in a higher temperature in the furnace freeboard. This requires greater amounts of cooling air to control the temperature. The furnace off-gas handling system capacity would have to be expanded to handle the greater quantities of gas. Also, the more metallized matte melts at higher temperatures, demanding superheated slags to control matte temperatures and bottom build-up. Refractory erosion in the slag zone with higher temperature slags must be controlled by cooling the refractory with copper coolers.
About 25% of the calcine escapes the six-in-line furnace; as much as possible of this is recycled back to the furnace20.
Inco""s roast-reduction smelting process25-29 involves deep roasting of nickel concentrate in fluidized-bed roasters. The roaster off-gas is treated in a sulphuric acid plant. The low-sulphur calcine is reduction smelted with coke in an electric furnace to yield a sulphur-deficient matte. This sulphur-deficient matte is converted to Bessemer matte in Peirce-Smith converters, with minimal evolution of sulphur dioxide (because of its sulphur-deficient nature), and the converter slag is returned to the electric furnace. Excellent recoveries of nickel were obtained, and the process was developed up to commercial-scale testing at the Thompson smelter during 1981 to 1982. Flash smelting of bulk copper-nickel concentrates was considered superior at Inco""s Copper Cliff smelter, but it was seen that in other circumstances the roast-reduction process could be an attractive option.
Sulphur is eliminated from the concentrate mainly in the roasters, running at 830 to 850xc2x0 C. The high temperatures promoted high oxygen efficiency, of approximately 95%. Slurry feeding permitted excellent control of the air to concentrate ratio in the roaster, and good control of sulphur elimination (approximately 80%). The process resulted in higher furnace temperatures, as well as higher iron levels to be oxidized in the converters.
U.S. Pat. No. 4,344,79225 describes the possibility of smelting either a partially roasted concentrate or a blend of dead-roasted concentrate and green concentrate, together with a carbonaceous reductant and silica flux. The feed is to contain only sufficient sulphur to produce a matte, in which the iron is present as metallic iron, and which has a sulphur deficiency of up to 25% with respect to the stoichiometric base metal sulphides Ni3S2, Cu2S, and Co9S8. The iron is later converted, to produce a low-iron matte by blowing and slagging the iron with silica flux, with very little release of sulphur dioxide during this stage of the process.
A process for the treatment of pyrrhotite, based on roasting to eliminate all or part of the sulphur, and hydrometallurgical treatment of the calcine to recover nickel, is described in Kerfoot30.
Sulphide ore concentrates containing platinum group metals (PGMs) have been roasted for various leaching processes.
The US Bureau of Mines devised a procedure for selectively extracting PGMs and gold from Stillwater Complex flotation concentrate. The concentrate was roasted at 1050xc2x0 C. to convert base-metal sulphides to oxides, and the PGMs from sulphide minerals to their elemental states. The roasted concentrate was then treated in a two-stage leaching process. Up to 97% of the platinum, 92% of the palladium, and 99% of the gold were extracted from the roasted concentrate31.
Other techniques are described in References 32 and 33.
Dead roasting of zinc concentrates is practised at industrial scale at Zincor, in Springs, South Africa. The calcine from this operation is treated by leaching and electrowinning.
A sulphide concentrate comprising 15% copper, 17% zinc, and 10% lead was roasted in a laboratory-scale fluidized bed in China, with the intention of using the product for further hydrometallurgical or direct smelting processing34.
Prime Western grade zinc has been produced from lead blast-furnace slags (and other zinc-containing waste materials) at large pilot-plant scale at Mintek in Randburg, South Africa, using the Enviroplas process35. Feed materials are smelted in a DC arc furnace, and the zinc is fumed off as a vapour, leaving behind a slag containing only small quantities of zinc oxide. The zinc vapour is subsequently treated in a lead splash condenser, resulting in the production of Prime Western grade zinc.
The invention provides a process for treating a metal sulphide concentrate which includes the steps of:
a) roasting the concentrate to reduce the sulphide content of the concentrate, and
b) smelting the concentrate, under reducing conditions, in an electrically stabilized open-arc furnace.
As used herein the phrase xe2x80x9can electrically stabilized open-arc furnacexe2x80x9d means a DC arc furnace or an electrically stabilized single electrode open-arc AC furnace (SOA furnace).
Preferably the roasting reduces the sulphide content to less than 10% sulphur by mass and, more preferably, to less than 1% of the initial amount present, the objective being to reduce the sulphide content to a negligible or otherwise acceptable value. The material produced by this step is referred to herein as xe2x80x9chighly-roastedxe2x80x9d or xe2x80x9cdead-roastedxe2x80x9d.
Preferably the roasting is done in a way which produces a steady stream of SO2-bearing gas. This gas may be used as a feedstock in a sulphuric acid plant. This step may be implemented in any appropriate way and for example the roasting may be performed in an enclosed vessel such as a fluidized bed reactor, to provide a high-concentration of SO2 in the gas.
Alternatively the SO2-bearing gas which is released in the roasting process may be subjected to gas scrubbing and neutralization.
The elimination of sulphur results in the valuable metals being collected in an alloy or from a vapour rather than as a matte, during the following smelting stage. This is believed to be advantageous as alloys have a greater collection efficiency than mattes.
The term xe2x80x9calloyxe2x80x9d is used here to denote a mixture of metals which may or may not contain some sulphur, as distinct from xe2x80x9cmattexe2x80x9d which is a mixture of metal sulphide.
The aforementioned process may be varied according to requirement and, more particularly, according to the nature of the metal sulphide or metal sulphide which are being treated.
When used for the treatment of zinc sulphide the calcine from the fluidized bed reactor may, optionally, be agglomerated before being fed to the arc furnace.
The reductant used in the furnace may be of any suitable kind and may for example be coke.
Zinc in the calcine is reduced to metal vapour and may be fumed off in a gas stream for recovery in any suitable way, for example in a lead splash condenser.
The aforementioned process as used for the treatment of zinc sulphide is particularly suitable for the treatment of zinc concentrates and zinc ore which contain relatively high levels of manganese, for example as encountered in the Gamsberg deposit in South Africa which has a manganese content which is higher by a factor of about 10 than the manganese content normally encountered in zinc concentrates.
The process may also be used for the treatment of nickel, copper and cobalt sulphide concentrates, whether existing as separate or combined sulphide, and PGM concentrates. The PGM concentrates may also be in the form of green furnace matte which is produced in any appropriate way, for example by making use of a six-in-line electric furnace.
The smelting of the dead-roasted concentrate, in the arc furnace, produces a slag which is depleted in metal values and an alloy. Iron in the alloy may optionally be removed, in oxide form, from the alloy using any suitable technique, such as by making use of a Peirce-Smith converter or a Top Blown Rotary Converter (TBRC).
Other undesirable elements such as carbon, silicon, or chromium, may be removed from the alloy using any suitable technique such as converting, or refining in a ladle furnace for example.
If the converter is used then alloy from the converter or, otherwise, alloy drawn directly from the arc furnace, if the converter is not used, is atomized so that it is in a form which is suitable for subsequent hydrometallurgical recovery of metal values, e.g. using a suitable leaching process.
Atomization of the alloy solves the problem of having to crush and mill an extremely tough alloy.
The green furnace matte may be granulated and milled, or water atomized, prior to the roasting step.
The smelting step may be a two-stage reduction smelting process, particularly when treating concentrates containing appreciable quantities of nickel and copper.
In the first stage, use could optionally be made of an arc furnace which is operated under slightly reducing conditions. This stage allows for the settling of some of the copper and nickel in an alloy, and a large fraction of the PGMs which partitions to the alloy.
In the second stage, which may be carried out in an arc furnace which is operated under highly reducing conditions, substantially all of the remaining nickel, the remaining PGMs, and most of the cobalt, are removed in an iron-based alloy which may also contain some copper.
The iron-based alloy may be atomized in preparation for hydrometallurgical treatment, e.g. for treatment in a leaching step.
The copper/nickel alloy from the first stage may be prepared for hydrometallurgical recovery by being water or gas atomized, granulated, or crushed and milled.
In carrying out the process of the invention use is made of a stabilised open arc furnace. A suitable furnace of this type is a DC arc furnace. The invention is however not limited in this regard for it may be possible to stabilise the arc or arcs of an AC arc furnace, using suitable control techniques, to achieve characteristics which are similar to those of a DC arc furnace in that the arc, or each arc, extends vertically from an overhead electrode to the charge, is confined, and does not deflect to side walls of the furnace.
A stabilized open arc furnace offers a number of advantages in the smelting of roasted sulphide concentrates and is seen as the enabling technology to make possible the process of the invention.
A DC arc furnace is roughly cylindrical in shape, often having a conical roof. A single vertical graphite electrode is used as the cathode, and the anode is embedded in the bottom of the furnace, in contact with the molten bath. The usual configuration involves operation with an open transferred plasma-arc above a molten bath with a surface substantially uncovered by feed materials (ie. an xe2x80x9copen bathxe2x80x9d operation). However, work has also been done using a two-electrode configuration (a twin-cathode is sometimes used for steel scrap or DRI (direct reduced iron) melting, and a two-electrode cathode-anode arrangement has also been used on a pilot scale.) Feed materials are either fed through the center of the electrode, or through a feed port fairly close to the electrode. Fewer feed ports are required with this configuration of furnace than are normally required for an AC six-in-line or a three phase three electrode AC furnace.
The powerful concentrated plasma arc jet provides a very efficient form of energy transfer to the molten bath of the furnace. This enables reactions to take place fairly rapidly, and good mixing is established in the bath, leading to a fairly uniform temperature distribution. The DC arc is relatively stable, not too easily extinguished, and is directed downwards towards the molten bath, with little flare towards the furnace walls. The arc jet xe2x80x98pullsxe2x80x99, via the Maecker effect, the furnace gases towards it, thereby attracting fine feed materials downwards into the bath, in so doing minimizing dust losses from the furnace. The low gas volumes from an electric furnace (compared to a furnace where energy is provided by combustion) also help in minimizing dust losses. The DC arc furnace can handle fine feed materials, typically sized below 3 mm, which makes it well suited for coupling to a fluidized-bed roaster.
The simple configuration of the DC arc furnace allows the freeboard to be well sealed, maintaining the CO atmosphere intemally, and minimizing the ingress of air.
Very high operating temperatures (much higher than those usually encountered in conventional base metals smelting) can be attained in the furnace, if required by the process, as power is supplied by the open arc, not merely by resistance heating in the slag. The furnace roof and walls are cooled (for example, by water-cooled copper panels) to retain the integrity of the furnace, even under conditions of high-intenisity smelting. Sigh freeboard temperatures are easily accommodated. The possibility of strongly reducing conditions in the furnace (together with the high operating temperature) avoids the common difficulties with the build-up of high-melting magnetite leading to operational problems in the furnace.
The processes described herein have high recoveries of the desired metals, and produce very clean slags. The DC arc furnace works well using iron alloy collection of valuable metals, or fuming off volatile metals. The processes result in low levels of impurities in the desired products.
The application of a DC or SOA arc furnace provides unique advantages particularly for feeds that contain high amounts of iron oxide which requires lots of reduction and for feeds that contain, for example, nickel and cobalt which require low oxygen potentials to achieve low slag losses.
A comparison of the characteristics of conventional furnaces and stabilised open arc furnaces highlights the advantages of using stabilised open arc furnaces in the process of the invention.
A conventional furnace has limitations in handling CO gas in the freeboard. Sealing the furnace is very difficult with multiple electrodes and feed points and a large cavity for the off-gas system. Rather than attempting to seal the furnace, the standard design involves the addition of air to combust the CO in the furnace freeboard and the addition of even more air to temper the combustion product gases. This results in large off-gas volumes, large quantities of dust make and the need to operate fans in a dirty gas environment.
In reduction smelting, reductants such as coke are mixed into the calcine. The reduction reaction is relatively slow in a conventional smelting configuration where the calcined material is smelted on a slag bath surface. The power density of the furnace which corresponds to the smelting rate cannot exceed the reduction reaction rate. This limitation becomes more important as the degree of roast increases and as the amount of reduction increases.
A calcine bed resting on the slag, or material banked up on the side walls, can roll over into the slag bath and cause unwanted slag foaming.
A conventional furnace (a rectangular six-in-line furnace, for example) requires good distribution of feed over the surface of the slag bath. This requires numerous feed points and a complex feed system above the furnace.
The reduction reaction is dependent on the reductant type. Generally, the need is for fine coke to maximize the reduction reaction rate. Otherwise, coke accumulates at the slag-calcine interface and redirects the furnace power, undesirably.
The metallized matte or metal that results from reduction smelting has a higher liquidus temperature. This necessitates a higher matte or metal temperature. A conventional furnace has poor capability to transfer energy in the vertical direction between the slag and matte phases.
A DC or SOA arc furnace:
(a) is small and intense;
(b) has no obvious limit on coke reduction kinetics. The ultimate case of dead roasting and back reduction to alloy is easily possible;
(c) can use a wide range of reductants eg. coal or coke of various size ranges;
(d) is easy to seal to contain a CO atmosphere, has little offgas, little dust, and few feed points; and
(e) produces hot metal
Recent developments in electrical power supply equipment have resulted in the possibility of using a three phase AC system to provide electrical energy via a single (graphite) electrode. This can be achieved by the switching of the three phase supply, possibly using pulse width modulation techniques, to generate a high frequency synchronized AC output. This development implies that a furnace configured in a similar manner to a DC arc furnace can be designed and used for similar process applications. Robicon of the USA have a power system (The Harmony Series) that can provide AC power as described above as well as the usual DC power output.
The stability of the AC high frequency arc is claimed to be better than a DC arc. There are a few possible disadvantages of the AC system:
the graphite electrode current capability may be less (skin effect);
the electrode wear may well be somewhat higher than with a DC furnace;
the arc jet and hence heat transfer to the bath may be slightly lower than for a DC arc furnace; and
the high frequency may generate harmonics although with suitable solid state switching techniques the harmonics may be reduced.
Potential benefits of an AC single electrode high frequency arc furnace include:
an improved arc stability (the arc is less likely to extinguish under certain circumstances); and
the arc may be less susceptible to loss of vertical directionality (e.g. sideways deflection) due to magnetic effects (deflection of the arc can damage the furnace sidewall and increase the energy losses).
Thus a suitably controlled power system can generate a high frequency waveform derived from a 3 phase alternating power supply which can be directly impressed across a single graphite electrode and a charge in a furnace, to produce a single open arc which is analogous to an arc in a DC furnace. This arrangement can offer similar and, in some regards improved, characteristics compared to a DC furnace and the scope of the invention therefore extends to the use of a DC arc furnace or a stabilised open arc AC furnace in the reduction smelting step.